1. Field of the Invention
This invention relates generally to the assay and recovery of precious metals.
More specifically, this invention relates to a method for the extraction of precious metals from complex ores and to their assay.
In one specific embodiment, this invention relates to the extraction of gold and platinum group metals from non-sulfidic placer ores and their concentrates.
2. Description of the Related Art
Fire assay has been one of the most widely used and most highly trusted techniques for the assay of ores and concentrates for gold, silver and the platinum group metals. In theory, the fire assay for precious metals depends primarily upon the solubility of the precious metals in molten metallic lead and upon their insolubility in slags of suitable composition. A complete separation of molten lead containing precious metals from slag is easily accomplished because of the large difference in specific gravity between the two liquids.
The fire assay, as it is typically carried out, is a two step process; the first step being a fusion and the second being cupellation. In the fusion step of the assay, a quantity of the sample, usually an ore or ore concentrate, is pulverized and mixed in a crucible with litharge, a reducing agent, and suitable fluxes so that the mixture will fuse at an easily attained temperature. Litharge, which is lead monoxide, is the source of metallic lead which serves to alloy with the precious metals. It also acts as a readily fusible basic flux. The reducing agent is typically flour or corn meal and serves to reduce a part of the litharge to metallic lead. It also functions to reduce certain metal oxides in the sample, particularly iron oxides, from a higher to a lower state. The fluxes used are selected to combine with the gangue portion of the ore sample to produce a relatively low melting point slag and so depend upon the composition of the gangue constituents. If the gangue is composed primarily of silica, which is an acid mineral, then basic fluxes such as lime, magnesia, sodium hydroxide, potassium carbonate, and others are employed. If, on the other hand, the gangue minerals are basic as are the lower oxides of iron and the oxides of manganese, sodium and potassium, then an acid flux such as silica or borax is used. Other fluxing agents are often added to modify the slag properties. Fluorspar, for example, increases the fluidity of most slags.
As the temperature of the charge is raised, the reducing agent reacts with part of the litharge to form molten metallic lead. That reaction begins at a temperature of about 500.degree. to 550.degree. C. Borax begins to melt at about the same temperature and forms fusible, glassy compounds with basic minerals of the ore charge and with basic flux constituents. At a somewhat higher temperature, on the order of 700.degree. C., lead oxide and silica start to combine and a viscous slag is formed. It is desirable to hold the charge near the slag formation temperature for a period of time long enough to decompose the ore particles and to allow the globules of metallic lead to alloy with the precious metals. Thereafter, the temperature is raised to a point where the slag is no longer viscous but is thin and fluid so as to allow ready settling of the lead globules to the bottom of the crucible. The maximum temperature attained during the fusion is ordinarily no higher than about 1,100.degree. C. At the end of the fusion, the molten mixture of metallic lead and slag is poured into a conical mold and is allowed to cool. A lead button, containing the precious metals, forms at the bottom of the mold and is readily separated from the glassy slag which contains the fluxes and the balance of the ore sample.
After separation from the slag, the lead button is subjected to a process called cupellation to separate the precious metals from the lead. Cupellation consists of an oxidizing fusion in a shallow, porous dish called a cupel, usually made of bone ash, which will absorb molten litharge but is impermeable to molten lead. The lead is oxidized at a temperature above the melting point of lead oxide which is mostly absorbed by the cupel and partly volatilized. At the end of the cupellation, after the lead has been driven off, there remains on the cupel a bead containing the precious metals. The bead is weighed and silver contained in the bead is parted from the gold and platinum metals using dilute nitric acid. The metal remaining is annealed and is then analyzed by various techniques to determine the proportions of the precious metals.
It is coming to be recognized that there exists many ores which are very difficult to assay with any degree of confidence using fire assay techniques or by chemical digestion using aqua regia and other powerfully oxidizing reagents followed by gravimetric or instrumental analysis of the extract. Such ores have been referred to by a variety of names including refractory ores, complex ores, unconventional ores, nonassayable ores and the like. Further, ores which are difficult to assay are almost always equally difficult to process for the recovery of contained precious metal values.
One approach to the recovery of precious metals from complex ores is set out in U.S. Pat. No. 4,892,631 to Merwin White. The process disclosed in that patent comprises the fusion of an ore in an electric arc furnace with fluxes and a base metal collector to which an precious metal inquart has been added. The base metal preferably is copper and the precious metal inquart is silver. Fluxes are used to produce an essentially neutral slag containing sodium, calcium, iron and aluminum gangue components of the ore as silicates. Precious metals released by the fusion transfer into the silver-inquarted copper. That copper product is then used as the anode of an electrolytic cell to plate the copper upon a cathode and to leave an anode residue comprising the precious metals. Copper plated upon the cathode may be recycled as the collector in a succeeding furnace charge.
The patent states that the optimum ratio by weight of the ore being smelted to the copper collector metal used is about 1:1. At that ratio, operation of the process would require that a ton of copper is subjected to electrolytic refining for each ton of ore smelted representing a considerable cost to the process.
It is also known to use other collector metals in the smelting of non-magnetic flotation or gravity concentrates containing platinum group metals. One such process is described in published South African Patent Application No. 78/2907 entitled "Process for Extraction of Platinum Group Metals from Chromite Bearing Ore". That process utilizes a high intensity wet magnetic separation to remove magnetic chromite from the non-magnetic platinum metal concentrate to thereby avoid reaction of chromite with carbon during the smelting step to form ferrochrome which alloys with platinum group metals. The smelting step itself comprises the formation of a furnace charge containing the non-magnetic platinum-containing concentrate together with collector materials for the platinum group metals, activators to improve the collection efficiency, and appropriate fluxes. The charge is smelted in a high intensity heating furnace, for example a plasma arc furnace, to form a slag layer and a layer containing collector material and platinum group metals. Collector material useful in the process comprises metals including copper, nickel, cobalt, iron, lead, zinc and mixtures thereof. The collector metals in an amount between about 3% and 10% by weight of the concentrates are added to the charge in metallic form or as the oxide or hydroxide which then undergoes reduction during the smelting step.